Method for recovering indium from indium-containing material

ABSTRACT

The present invention provides a recovery method in which the indium sponge that is deposited by substitution in the substitution deposition step in the recovery of indium from an indium-containing material is produced in the form of a powder rather than in a bulky form. This recovery method is a method in which an indium sponge is deposited by substitution by adding a substance containing chlorine, and further adding a reducing agent, to an indium-containing solution whose pH has been adjusted to a value in the range of 1 to 2.2.

BACKGROUND OF THE INVENTION

1. Field of the Invention

The present invention relates to a method for recovering indium from anindium-containing material.

2. Description of the Related Art

As a conductive compound of the III-V group, indium is utilized inintermetallic compounds such as InP, InAs, and the like, in indium oxidedoped with tin (ITO) as a material for use in solar cells, and intransparent conductive thin films. It is expected that the demand forindium will become increasingly greater in the future.

Conventionally, there has been no principal ore for indium;industrially, indium has been produced by recovering indium as aby-product in zinc purification or lead purification, e.g., byrecovering indium concentrated in soot and smoke. Accordingly, rawmaterials used for the recovery of indium contain large amounts of metalimpurities such as Zn, Fe, Cu, Al, Ga, As, Cd, and the like.Furthermore, many types of components other than these metal componentsare also contained in trace amounts.

Accordingly, a complicated process is required in order to remove thesemetal impurities and recover high-purity indium. Generally, theabovementioned indium recovery process is accomplished by a combinationof electrolytic purification methods and chemical purification such as(A) methods in which indium is precipitated as a hydroxide by adjustingthe pH, (B) methods in which indium is precipitated as a sulfide byadding a sulfurizing agent, (C) methods in which indium is substitutedand deposited by adding metals such as Al, Zn, Cd, Zn—Cd alloys, and thelike, (D) methods in which indium is recovered by solvent extraction,(E) methods in which indium is recovered by an ion exchange method, andthe like.

However, in the abovementioned recovery processes, methods according to(A) are methods which utilize differences in the pH regions where metalions produce hydroxides. For example, a method in which Zn and Al aredissolved and In is precipitated and recovered as a hydroxide by raisingthe pH to 12 or greater is used as a method for separating In from Znand Al. In this method, however, the In hydroxide that is produced hasextremely poor filtration characteristics; accordingly, the size of thefiltration equipment is increased, and the operation also requires along time. In this method, it is also difficult to separate In fromimpurities such as Fe, Cu, As, Cd, and the like.

The methods of (B) utilize differences in the solubility product ofmetal sulfides: in such methods, low-purity sulfides containing variousmetal impurities such as those described above are produced in largequantities. These sulfides generally have poor filtrationcharacteristics; furthermore, in cases where the In sulfide obtained isextracted, it is difficult to extract the In completely using sulfuricacid alone. Accordingly, such methods suffer from the problem of beingdifficult to apply to wet zinc processes.

A problem is involved in (C) such that in cases where impurities thatare nobler than In are contained, separation of In from these metals isimpossible.

A problem is involved in (D) and (E) such that a burden is placed on thepre-processing by the impurities that are separated from In;furthermore, the running cost is high.

In all of the chemical purification methods described above, theseparation of impurity metals is inadequate. Accordingly, neither can asimple electrowinning method (in which the object metal is extractedinto an aqueous solution, electrolysis is performed using an insolubleanode, and a high-purity metal is obtained in a single process at thecathode) be used as a combination electrolytic purification method, andit has been unavoidably necessary to use a cumbersome electrolyticpurification method (in which crude metal is used as the anode, andpurification is performed by electrolysis of the high-purity metal atthe cathode).

Accordingly, each of the methods described above has a respectiveproblem. Combinations of the abovementioned methods are used for actualrecovery, and the process used in order to recover high-purity In becomecomplicated and bothersome; this has not been an economical method.

In Patent Document 1, the present inventors proposed a method forrecovering indium from an indium-containing material including an acidextraction step in which the indium-containing material is subjected toextraction processing with an acid, and metals that are soluble in theacid are dissolved together with In; a step of removing Cu and the likein which a sulfurizing agent is added to the extract obtained in theabovementioned acid extraction step while adjusting theoxidation-reduction potential, and metals other than In such as Cu andthe like are removed by precipitation; a sulfurizing and precipitationstep in which sulfuric acid and a sulfurizing agent are added to theaqueous solution of indium obtained in the abovementioned step ofremoving Cu and the like, and In is precipitated and concentrated as asulfide; an SO₂ extraction step in which In is selectively extracted byblowing SO₂ gas into the indium sulfide obtained in the abovementionedsulfurizing and precipitation step in the presence of sulfuric acidacidity; a substitution deposition step in which the pH and solute SO₂concentration of the indium-containing extract obtained in theabove-mentioned SO₂ extraction step is adjusted, a metal powder is thenadded, and an indium sponge is deposited by substitution; a hydrochloricacid extraction step in which the indium sponge obtained in theabovementioned substitution deposition step is extracted withhydrochloric acid; a step of removing Cd and the like in which asulfurizing agent is added to the indium extract obtained in theabovementioned hydrochloric acid extraction step, residual metal ions ofCd and the like are removed by precipitation, and a startingelectrolytic solution is obtained; and an electrowinning step in whichthe starting electrolytic solution obtained in the abovementioned stepof removing Cd and the like is electrolyzed, and high-purity metallicindium is obtained.

[Patent Document 1] Japanese Laid-Open Patent Application No. 11-269570.

SUMMARY OF THE INVENTION

The steps of recovering indium from an indium-containing material havebeen simplified by the method described in Patent Document 1, and thismethod has contributed to a reduction in the production cost. However,using the method described in Patent Document 1 may involve the problemsuch that the indium sponge that is deposited by substitution in thesubstitution deposition step is produced as a bulky form. When a bulkyindium sponge is produced, the hydrochloric acid extraction cannot beperformed unmodified in the hydrochloric acid extraction step.Accordingly, a step in which this bulky indium sponge is pulverized isnewly required, and this pulverization step has been the cause of anincrease in the production cost.

The present invention was devised in light of the abovementionedconditions; it is an object of the present invention to provide a methodfor recovering indium from an indium-containing material in which theindium sponge deposited by substitution in the substitution depositionstep is produced not in bulk form, but rather in powder form.

MEANS USED TO SOLVE THE PROBLEMS

The present inventors conducted diligent research in order to solve theabovementioned problem. As a result of trial and error, the inventorsdiscovered that the conversion of the indium sponge into a bulky formcan be suppressed by performing a step of adding a substance containingchlorine to the indium-containing solution whose pH has been adjusted toa value in the range of 1 to 2.2 in the above-mentioned substitutiondeposition step, and a step of adding a reducing agent to theindium-containing solution to which the abovementioned substancecontaining chlorine has been added, and depositing an indium sponge bysubstitution. The inventors perfected the present invention as a resultof this discovery.

Specifically, in order to solve the aforementioned problem, a firstaspect provides a method for recovering indium from an indium-containingmaterial, wherein this method has a step of adding a substancecontaining chlorine to an indium-containing solution whose pH has beenadjusted to a value in the range of 1 to 2.2, and a step of adding areducing agent to the indium-containing solution to which theabovementioned substance containing chlorine has been added, anddepositing an indium sponge by substitution.

A second aspect of the present invention provides the method forrecovering indium from an indium-containing material according to thefirst aspect, wherein the value of the molar ratio of Cl/In is set at avalue that is greater than 0 but no greater than 0.68 in theabovementioned step in which a substance containing chlorine is added.

A third aspect of the present invention provides the method forrecovering indium from an indium-containing material according to thefirst or second aspect, wherein one or more substances selected from thegroup consisting of sodium chloride and an electrolytic tailing solutionof indium is used as the abovementioned substance containing chlorine.

A fourth aspect provides the method for recovering indium from anindium-containing material according to any of the first through thirdaspects, wherein metallic zinc is used as the abovementioned reducingagent.

EFFECT OF THE INVENTION

In these aspects of the present invention, the indium sponge depositedby substitution in the substitution deposition step is produced aspowder, and the productivity of the recovery of indium from anindium-containing material is improved.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a process diagram showing an outline of the method of thepresent invention.

DESCRIPTION OF THE PREFERRED EMBODIMENTS

In the present invention, substances containing indium can be widelyused as starting raw materials; here, however, a case in which thepresent invention is applied to neutral gypsum which is produced as aby-product in wet zinc purification will be described as an example. Anexample of the indium recovery process by this method is shown in FIG.1.

In step (1), when neutral gypsum is extracted using sulfuric acid,impurity metal ions such as Cu, As, Al, Fe, Zn, Ga, and the like thatare soluble in acids are extracted together with In, and a slurry withinsoluble gypsum is formed. Besides sulfuric acid, it would also bepossible to use hydrochloric acid, nitric acid, or the like as the acidused in this extraction. The present invention is not limited tosulfuric acid; however, sulfuric acid is the least expensive. Thesulfuric acid concentration of the In extract is ordinarily 20 to 40g/l.

In step (2), for example, H₂S or NaSH is added as a sulfurizing agent tothe In extract slurry obtained in step (1) while the oxidation-reductionpotential (hereafter referred to as “Eh”) is controlled so that thispotential is in the range of 50 to 320 mV (using an Ag/AgCl electrode),and impurities such as Cu, As, and the like are precipitated and removedas sulfides. In this case, since the sulfuric acid concentration is alsocontrolled to 20 to 40 g/l, In is not precipitated.

90% or more of the In contained in the neutral gypsum is transferredinto the acidic solution of sulfuric acid by the processing in steps (1)and (2). Accordingly, for example, the precipitate (copper residue) issubjected to a solid-liquid separation using a filter press or the like,In this case, the insoluble gypsum acts as a filtration assistant duringextraction; accordingly, the generally poor filtration characteristicsof sulfides are greatly improved. The copper residue is sent back intothe main system of zinc purification.

In step (3), a sulfurizing agent, e.g., H₂S or NaSH is addedsimultaneously with sulfuric acid to the aqueous solution containing Inobtained in step (2), and In is precipitated as a sulfide. This issubjected to a solid-liquid separation using a filter press or the like,and impurities such as Zn, Fe, Al, Ga, and the like remaining in thesolution are separated and removed. The recovery rate of In in theprecipitate is 95% or greater. The filtrate (sulfurized solution) issent to the drainage system.

In step (4), In is extracted while SO₂ gas is blown into the indiumsulfide obtained in step (3) under sulfuric acid acidity.

Acid extraction methods for sulfides generally include three types,i.e., (a) the hydrogen sulfide generating type, (b) the sulfur producingtype, and (c) the sulfuric acid producing type. However, in cases whereindium sulfide is extracted, the solubility product in the reactionoperating arm (a) is small, and therefore In cannot be completelyextracted using this method. In cases where oxygen is used as anoxidizing agent in the reactions of (b) and (c), the reactiontemperature and pressure must be respectively set at high values, i.e.,150° C. and 12 kg/cm²; accordingly, a pressure vessel such as anautoclave or the like must be used as the reaction vessel. Furthermore,although In can be completely extracted in these methods, the oxidizingpower is strong, so that impurities that are contained are alsosimilarly completely extracted.

In the method of the present invention, a combination of the reactionsof (a) and (b) is performed by using SO₂ as an oxidizing agent, theoxidizing power is appropriately controlled, and the extraction of otherimpurities is suppressed while In is extracted; i.e., In is selectivelyextracted. The temperature in this case may be ordinary temperature, andthe pressure may also be atmospheric pressure. Accordingly, an ordinaryreaction vessel can be used. After the reaction, 90% or more of the Inis transferred into the extract; accordingly, a solid-liquid separationis performed using a filter press or the like. The cake (sulfur residue)is sent to the main system of zinc purification.

In step (5), the In extract obtained in step (4) is neutralized with analkali, e.g., caustic soda or the like, and the pH is preferablyadjusted to a value in the range of 1 to 2.2. The associated rationaleis that if the pH is 1 or greater, the amount of zinc powder that isadded as a substitution agent in a later step can be suppressed, and ifthe pH is 2.2 or less, the bulk formation of the In sponge deposited ina later step can be suppressed. Following the adjustment of the pH, asubstance containing chlorine (e.g., a chlorine compound that is a saltof sodium, or an electrolytic tailing solution of indium) is added, andthe value of the molar ratio of Cl/In in the In extraction is adjustedto a value of 0 to 1.44, preferably 0.68 or less. Next, a powder of ametal which has a greater tendency toward ionization than indium, e.g.,a zinc powder, is added, and a powder-form indium sponge is deposited bysubstitution It is desirable that the oxidation-reduction potential be−200 mV or less.

Since SO₂ is used in the extraction in step (4), SO₂ is dissolved in theIn extraction that is supplied to step (5). Accordingly, the bulkyindium sponge can be prevented from forming, and a powder-form indiumsponge can be obtained, by blowing air in, and controlling the SO₂concentration in the In extract to a value of 0.05 to 0.3 g/l. Thesubstituted solution is returned to the abovementioned step (3).

In step (6), the powder-form indium sponge obtained in step (5) isextracted using hydrochloric acid, with the pH controlled to a value inthe range of 0.5 to 1.5 and the Eh controlled to a value in the range of−400 to −500 mV. In this case, 90% or more of the In is transferred intothe extract; accordingly, a solid-liquid separation is performed using afilter press or the like. Trace metals such as Cd, Pb, Ni, As, and thelike can be concentrated and removed in the extraction residue (spongeresidue). The sponge residue is returned to the abovementioned step (4).

In step (7), in cases where Cd, As, and the like still remain in the Inextract obtained in step (6), a sulfurizing agent, e.g., H₂S gas, isblown in, a final cleaning is performed, a solid-liquid separation isperformed, and the filtrate is used as a starting electrolytic solution.The cake (cadmium residue) is returned to the abovementioned step (4).

In step (8), electrowinning is performed using a DSA (dimensionallysuitable anode) as the anode, and a Ti plate as the cathode, from thestarting electrolytic solution obtained in step (7), and high-paritymetallic indium is obtained.

EXAMPLES

The recovery processing of indium was performed using neutral gypsumproduced as a by-product in a wet zinc purification process as thestarting raw material. (1) Acid extraction: Water was added to 294.5 gof neutral gypsum which was the raw material for In recovery, to form apulp having a solid concentration of 203 g/l. While mechanical agitationwas performed using an agitator, sulfuric acid was added to this so thatthe final acid concentration was 28 g/l, and an extraction was performedfor 2 hours while the temperature was maintained at 60° C. The contentsand distributions of In, Zn, Cu, and As in the raw material and theextract obtained are shown in Table 1. TABLE 1 Material balance of acidextraction step Amount Content (%, g/l) Distribution (%) (g, ml) In ZnCu As In Zn Cu As Raw material 294.5 1.54 13.1 3.49 1.28 100.0 100.0100.0 100.0 Extract 1450 3.03 25.8 6.86 2.47 96.7 97.1 96.8 94.9

In Table 1, the raw material amounts are shown in g, and the amounts inthe extract are shown in ml.

(2) Removal of Cu and the like: NaSH was added to the extract slurryobtained in the abovementioned extraction step, and a sulfurizingreaction was performed until Eh reached 300 mV (using an Ag/AgClelectrode). The reaction time was 2 hours, and the reaction temperaturewas 60° C. Following the completion of the reaction, the slurry obtainedwas filtered; the cake was used as a copper residue, and the filtratewas used as a copper-free solution. The respective analysis results areshown in Table 2. TABLE 2 Material balance of step of removing Cu andthe like Amount Content (%, g/l) Distribution (%) (g, ml) In Zn Cu As InZn Cu As Extract slurry 1500 3.20 27.4 7.26 2.61 100.0 100.0 100.0 100.0Copper residue 107.1 0.13 0.30 10.2 1.70 2.9 0.8 100.3 46.5 Copper-free1450 3.21 28.1 tr 1.44 97.1 99.2 0.0 53.5 solution

In Table 2, the amounts in the extract slurry and the amounts in thecopper-free solution are shown in ml, and the amounts in the copperresidue are shown in g.

(3) Sulfurizing and precipitation: While the abovementioned copper-freesolution (aqueous solution containing In) was agitated with an agitator,the pH was maintained at a constant level of 0.8 with sulfuric acid;NaSH was added until Eh reached −20 mV (using an Ag/AgCl electrode), andIn was precipitated as a sulfide. The reaction was performed for 5 hoursat a temperature of 60° C. Following the completion of the reaction, theslurry obtained was filtered; the cake was used as a sulfurizationresidue, and the filtrate was used as a sulfurized solution. Therespective analysis results and material balances are shown in Table 3.TABLE 3 Material balance of sulfurization precipitation step AmountContent (%, g/l) Distribution (%) (g, ml) In Zn Cu As In Zn AsCopper-free 1450 3.21 28.1 tr 1.44 100.0 100.0 100.0 solution Sulfurized1430 0.02 24.6 0.0 tr 0.6 86.3 0.0 solution Sulfurization 45.7 9.96 22.80.0 4.55 97.8 25.6 99.6 residue

In Table 3, the amounts in the copper-free solution and the amounts inthe sulfurized solution are in ml, and the amounts in the sulfurizationresidue are in g.

(4) SO₂ extraction: The sulfurization residues obtained by repeating theabovementioned steps (1) through (3) were collected to produce 417.7 g;water was added to this to form pulp with a solid concentration of 119q/l. While this was agitated with an agitator, sulfuric acid was addedso that the sulfuric acid concentration was adjusted to 51 g/l, and SO₂gas was blown in so that the solute SO₂ concentration was 8 g/l. Thereaction was performed for 2 hours at a temperature of 80° C. Followingthe completion of the reaction, the slurry obtained was filtered, thecake was used as a sulfur residue, and the filtrate was used as an SO₂extract. The respective analysis results and material balances are shownin Table 4. TABLE 4 Material balance of SO₂ extraction step AmountContent (%, g/l) Distribution (%) (g, ml) In Zn Cu As In Zn AsSulfurization 417.7 13.41 19.1 0.0 10.32 100.0 100.0 100.0 residueSulfur residue 236.0 0.78 3.75 0.0 17.50 3.3 11.1 95.8 SO₂ extract 345018.0 30.7 0.0 0.20 110.9 132.6 1.6

In Table 4, the amounts in the sulfurization residue and the amounts inthe sulfur residue are in g, and the amounts in the SO₂ extract are inml.

(5) Substitution deposition: Air was blown into the abovementioned SO₂extract, and degassing was performed until the solute SO₂ concentrationwas 0.2 g/l. NaOH was added, and the extract was neutralized until thepH reached 2.2. This was used as a substitution raw-material solution.3000 ml of this substitution raw-material solution was taken as onesample, and three samples (samples 1 through 3) were thus prepared. Insample 1 thus prepared, sodium chloride was added, and the value ofCl/In molar ratio was adjusted to 0.68 or less. In sample 2, sodiumchloride was added, and the value of Cl/In molar ratio was adjusted to0.68 to 1.44, In sample 3, sodium chloride was added, and the value ofCl/In molar ratio was adjusted to 1.44 or greater.

1.8 equivalents of zinc powder with respect to the In was added to the3000 ml of each of the substitution raw-material solution in theprepared samples 1 through 3, and an In sponge was deposited bysubstitution. The reaction temperature was 60° C., and the reaction timewas 1 hour. The probability of bulky sponge generation in each sample isshown in Table 5, and the analysis results and material balance for eachproduct are shown in Table 6. TABLE 5 Probability of generation of bulkysponge in substitution process No. of times Probability of bulky formoccurrence of Cl/In molar Batch No. occurred bulky form Sample ratio(times) (times) (%) Sample 1 ≦0.68 73 0 0 Sample 2 0.68 to 1.44 47 14 30Sample 3 ≧1.44 9 9 100

TABLE 6 Material balance of substitution deposition step Amount Content(%, g/l) Distribution (%) (g, ml) In Zn Cu As In Zn As Starting 300018.0 30.7 tr 0.20 100.0 100.0 100.0 substitution solution Substituted2900 0.10 55.3 0.00 tr 0.5 174.1 0.0 solution In sponge 62.1 87.0 3.500.0 0.92 100.0 2.4 95.2

In Table 6, the amounts in the substitution raw-material solution andthe amounts in the substituted solution are in ml, and the amounts inthe In sponge are in g.

(6) Hydrochloric acid extracting: Water was added to 238.1 g of the Insponge collected by repeating the various steps described above, thusforming a pulp having a solid concentration of 144 g/l. While this wasagitated with an agitator, hydrochloric acid was added so that the pHwas adjusted to 1, and Eh to −480 mV (using an Ag/AgCl electrode), andindium was extracted. The reaction temperature was 65° C., and thereaction time was 3 hours. The analysis results and material balances ofthe respective products are shown in Table 7. TABLE 7 Material balanceof hydrochloric acid extraction step Amount Content (%, ppm, g/l, mg/l)Distribution (%) (g, ml) In Ni Cd Pb As In Ni Cd Pb As In sponge 238.187.9 122 4400 4400 1780 100 100 100 100 100 Sponge residue 47.1 46.52 6220 222 90 10.5 95.4 98.8 100.0 99.5 Hydrochloric 1500 130.6 0.9 8.7 0.20.2 93.6 4.9 1.3 0.0 0.1 acid extract

In Table 7, the amounts in the In sponge and the amounts In spongeresidue are in g, the amounts in the hydrochloric acid extract are inml, the In contents are in % and g/l, and the contents of componentsother than In are in ppm and mg/i.

(7) Step of removing Cd and the like: NaOH was added to 1500 ml of thehydrochloric acid extract obtained in the abovementioned hydrochloricacid extraction step, and the extract was neutralized to a pH of 1.5.1.5 l of H₂S gas was blown into this solution, and impurities such as Cdand the like were precipitated as sulfides. The reaction temperature was40° C., and the reaction time was 0.5 hours. The suspension followingthe reaction was filtered, the cake was used as a cadmium residue, andthe filtrate was used as a cadmium-free solution. The analysis resultsand material balances of the respective products are shown in Table 8.TABLE 8 Material balance of step of removal of Cd and the like AmountContent (%, ppm, g/l, mg/l) Distribution (%) (g, ml) In Ni Cd Pb As InNi Cd Pb As Hydrochloric acid 1500 130.6 0.9 8.7 0.2 0.2 100 100 100 100100 extract Cadmium residue 2.6 50.35 0.0 2.7 0.1 0.0 0.7 0.0 98.9 100.00.0 Cd-free solution 1500 129.5 1.2 0.1 0.0 0.2 99.2 133.3 1.1 0.0 100.0

In Table 8, the amounts in the hydrochloric acid and the amounts in theCd-free solution are in ml, the amounts in the Cd residue are in g, theIn contents are in % in g/l, and the contents of components other thanIn are in ppm and mg/l.

(8) Electrowinning step: The Cd-free solution obtained in theabovementioned step (7) was used as a starting electrolytic solution,and electrowinning was performed for 48 hours at a temperature of 40° C.and a current density of 150 A/m². A DSA was used for the anode, and aTi plate was used for the cathode. The analysis results and materialbalances of the starting electrolytic solution and the indium andelectrolytic tailing solution obtained are shown in Table 9. TABLE 9Material balance of electrowinning step Amount Content (%, ppm, g/l,mg/l) Distribution (%) (g, ml) In Ni Cd Pb As In Ni Cd Starting 280070.0 0.7 0.4 0.0 0.0 100.0 100.0 100.0 electrolytic solution Indium 15099.99994 0.3 0.0 0.2 0.0 76.5 2.3 0.0 Electrolytic 2800 16.4 0.7 0.4 0.00.0 23.4 100.0 100.0 tailing solution

In Table 9, the amounts in the starting electrolytic solution and theamounts in the electrolytic tailing solution are in ml, the amount ofindium is in g, the content of In is in % and g/l, and the contents ofcomponents other than In are in ppm and mg/l.

1. A method for recovering indium from an indium-containing material,comprising: a step of adding a substance containing chlorine to anindium-containing solution whose pH has been adjusted to a value in therange of 1 to 2.2; and a step of adding a reducing agent to theindium-containing solution to which said substance containing chlorinehas been added, and depositing an indium sponge by substitution.
 2. Themethod for recovering indium from an indium-containing materialaccording to claim 1, wherein the value of the molar ratio of Cl/In isadjusted to a value that is greater than 0 but no greater than 0.68 inthe step of adding said substance containing chlorine.
 3. The method forrecovering indium from an indium-containing material according to claim1, wherein one or more substances selected from the group consisting ofsodium chloride and an electrolytic tailing solution of indium are usedas said substance containing chlorine.
 4. The method for recoveringindium from an indium-containing material according to claim 2, whereinone or more substances selected from the group consisting of sodiumchloride and an electrolytic tailing solution of indium are used as saidsubstance containing chlorine.
 5. The method for recovering indium froman indium-containing material according to claim 1, wherein metalliczinc is used as said reducing agent.
 6. The method for recovering indiumfrom an indium-containing material according to claim 2, whereinmetallic zinc is used as said reducing agent.
 7. The method forrecovering indium from an indium-containing material according to claim3, wherein metallic zinc is used as said reducing agent.
 8. The methodfor recovering indium from an indium-containing material according toclaim 4, wherein metallic zinc is used as said reducing agent.